Low-cost selective precipitation circuit for recovery of rare earth elements from acid leachate of coal waste

ABSTRACT

The present invention concerns a process of selective precipitation for the purpose of recovering rare earth elements from acidic media derived from coal and coal byproducts via two main steps of sequential precipitation and selective precipitation. An intermediary step of re-precipitation can be included to further increase RRE concentrations, as well as improve contaminant metal removal.

RELATED APPLICATIONS

This application claims priority to U.S. Provisional Patent Application62/583,644, filed Nov. 9, 2017, all of which is hereby incorporated byreference in its entirety.

GOVERNMENT SUPPORT

The present invention was made with support from the US Department ofEnergy under grant DE-FE0027035. The Government may have certain rightsto the invention.

FIELD OF THE INVENTION

The present invention relates to a process of selective precipitationfor the purpose of recovering rare earth elements from acidic mediaderived from coal and coal byproducts.

BACKGROUND

Rare earth elements (REEs) were found to be leached out from coal coarserefuse piles under the effects of acids generated through pyriteoxidization which occurred naturally. The ratio of critical touncritical REEs of the leachate can be as high as 2.0 which issignificantly higher than the coarse refuse (about 0.25). The terms“rare earths” and “rare earth elements” as used herein include scandiumand yttrium, as well as those elements having atomic number from 57 to71 inclusively.

That rare earth elements occur in coal and coal byproducts is widelyknown and has been a topic of historic and ongoing research as shown bySeredin, V. V., & Dai, S. (Coal deposits as potential alternativesources for lanthanides and yttrium. International Journal of CoalGeology, 94, 67-93, 2012) and Honaker et al. (Process evaluation andflowsheet development for the recovery of rare earth elements from coaland associated byproducts. Minerals & Metallurgical Processing, 34(3),107-115, 2017).

However, it is not commonly known that, due to the nature and attributesof coal and coal byproducts, natural and synthetic leachates can and areformed containing rare earth elements. These leachates may be processedin such a way as to increase the concentration of rare earth elements asinsoluble precipitates.

A review of prior art shows that upgrading of rare earth elements viaprecipitation is not novel. Kaneyoshi et al. U.S. Pat. No. 5,545,386describe a method of producing rare earth elements via pH manipulationto produce a precipitate that is treated with heat to produce an oxide.This work assumes that the rare earth elements were initially watersoluble. The solution is then mixed with an alkaline solution toprecipitate hydroxides. The precipitate is then separated from theaqueous medium and calcined to produce an oxide.

Further, Fulford et al. U.S. Pat. No. 5,015,447 describe a sulfuricbased aqueous rare earth solution. This invention covers the utilizationof solvent extraction techniques to upgrade rare earth elements byconcentration via extractants specified in the organic phase. The rareearth elements are then extracted from the organic phase with increasedpurity and concentration. The operation of the organic extraction isachieved through pH manipulation.

SUMMARY OF THE INVENTION

The present invention describes a process of manipulating the pH of anacidic liquid containing earth elements in such a way as to selectivelyprecipitate rare earth elements from therein, thus separating the rareearths from contaminants which may be contained in the acidic liquid.The acidic liquid may be generated naturally via acidification of coaland coal byproducts by the oxidation of contained pyrite and by contactwith rain or ground water. Natural sources include all sources ofsufficient acidity as to contain rare earth elements from coal miningderived sources. In addition, synthetic solutions may be obtained fromleaching under controlled processing. These may include but are notexclusive to tank or heap leaching. Further, in the treatment of acidicwaters generated from the storage of coal sources, precipitates producedby pH manipulation by mine operators are created containing rare earthelements. These precipitates may also be considered a feed material,with the material being re-leached to bring the rare earth elementsagain into solution.

The present invention recovers rare earth elements by three steps: 1)sequential precipitation; 2) re-precipitation (optional); and 3)selective precipitation.

In the sequential step, the rare earth containing solution is pHmanipulated via controlled and incremental basic solution addition toprecipitate out various contaminative elements. The mode of solutionaddition and aging time allows for the precipitation of elements ofinterest. After removal of the contaminative element precipitates, thepH is slowly raised to precipitate rare earth elements.

The re-precipitation step is then used to further dissolve precipitatesformed in the sequential precipitation step and treat them in a mannerwhich further removes contamination. Thus, depending on the initialsolution, this step is optional. An example of use would be whenaluminum concentration in the feed is high.

The final step of selective precipitation is when the intermediate rareearth solution is re-filtered to remove insoluble precipitates followedby conditioning with an oxalic acid solution and then pH adjustment. Therare earth elements are selectively precipitated and recovered byfiltration with the exception of scandium which remains in the filtrate.Scandium recovery is achieved and a concentrate produced by treating thefiltrate by solvent extraction.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 shows a schematic overview of the steps described in the presentinvention, configured for feeds exhibiting low aluminum content whichincludes the obtaining of precipitate 1 by sequential precipitation; andthe selective precipitation (precipitate 3). The label numberscorrespond to processing steps defined by the embodiment examples.

FIG. 2 shows a schematic overview of the steps described in the presentinvention, including the obtaining of precipitate 1 by sequentialprecipitation; the optional re-precipitation step to obtain precipitate;and the selective precipitation (precipitate 3) from a re-leachedsolution in a different acid class (reagent 1). The label numberscorrespond to processing steps defined by the embodiment examples.

FIG. 3 shows REE concentration in acid mine drainage as a function ofsolution pH values based on data reported in literature. (The data arefrom Da Silva et al., 2009; Migaszewski et al., 2016; Sahoo et al.,2012; Stewart et al., 2017; Sun et al., 2012; Worrall and Pearson,2001a, Worrall and Pearson, 2001b; Zhao et al., 2007; Ziemkiewicz etal., 2016).

FIG. 4 shows Percentages of individual REEs in the total REEs for boththe coarse refuse and natural leachate samples.

FIG. 5 shows NASC normalized patterns of both the coarse refuse andnatural leachate samples of West Kentucky No. 13 seam (rare earthcontents of NASC were referred to Gromet et al., 1984).

FIG. 6 shows Recovery of total, critical and heavy REEs, scandium, andthorium in the precipitates obtained in different pH ranges (recoveryvalues were calculated on a mass basis and add to 100% for eachmetal/group).

FIG. 7 shows Solution pH changes of the natural leachate as a functionof the amount of sodium hydroxide added and appearance of theprecipitates obtained from the staged precipitation test.

FIG. 8 shows Saturation indexes of the minerals possibly formed in theleachate at different pH values: (a) rare earth hydroxides (amorphous);(b) copper and zinc minerals; (c) iron and (d) aluminum minerals.

FIG. 9 shows Percentages of total metals in form of free ions in thesolution (calculated using molar concentrations of different metalspecies).

FIG. 10 shows Removal of iron, aluminum and lanthanum from the modelsolutions as a function of the solution pH values: (a) iron andaluminum; (b) lanthanum.

DETAILED DESCRIPTION

The present invention relates to a process for enriching rare earthelements within a leachate precipitate. In the process, contaminantmetals, such as iron and aluminum, are first sequentially eliminatedfrom the leachate by raising the solution pH values to about 3.5 and4.5. The exact pH values for a specific leachate are determined byanalyzing the concentration changes of iron, aluminum and REEs withincreases in solution pH values. Neutralizing reagents such as sodiumhydroxide are added in micro increments. After each addition, thesolution is stirred for a few minutes to re-dissolve the precipitateswhich are formed due to inefficient dispersion of the neutralizingreagent. Instead of filtering out the precipitates until the solution pHvalues reach 3.5 or 4.5, iron and aluminum precipitates are removed aslong as measurable amounts are formed. This practice can reduce the lossof REEs to iron and aluminum precipitates due to surface adsorption.

As an overview, precipitates containing about 0.5-1% of REEs areobtained in the pH range about 5.0-9.0 through the sequentialprecipitation process. The majority of iron is removed while aluminummay still be the dominant contaminant due to its higher precipitationpoint and higher concentration in some leachate samples. To furtherreduce the aluminum content, the precipitate is optionally re-dissolvedin smaller volume of solution at pH 2.0. For example, precipitatesobtained from 10 liters of leachate may be re-dissolved in 1 liter ofsolution. In this case, aluminum can be removed at lower pH values(e.g., 3.5) due to its increased concentration, while a minimum amountof REEs is lost. The solution that is used to re-dissolve theprecipitates can be the residual leachate after the precipitate isfiltered out at pH 9.0. It is advisable that the dissolving acid in thisstep be different than that of the initial step so as to promoteadditional precipitation selectivity. More efficient removal of aluminumcan be obtained by adding inorganic regulators such as MgCl2 or CaCl2)into the solution. The purpose of these is to prevent REEs fromco-precipitating or adsorbing on the contaminant precipitates. REEs arere-precipitated from the solution in the pH range 5.0-9.0. It isadvisable that the dissolving acid in this step be different than thatof the proceeding step so as to promote additional precipitationselectivity.

The REEs-enriched precipitates obtained from the above steps arere-dissolved in smaller volumes of liquid (e.g., 100 ml versus a startleachate volume of 10 liters). It is advisable that the dissolving acidin this step be different than that of the initial step so as to promoteadditional precipitation selectivity. The REEs concentrations in theliquid are in the range of 0.1-0.2 g/L with major contaminants of Mg,Mn, and Ca. REEs are precipitated out from the liquid using a differentclass of acid, such as nitric and oxalic acid, in the pH range of1.0-1.2, which can generate a final rare earth product containing about70% or higher. The REEs remaining in solution containing primarilyscandium can also be further purified using a simple solvent extractionprocess. The process combines a single stage of loading, scrubbing andstripping. The REEs in the stripping solution are finally precipitatedout using oxalic acid in the pH range of 1.0-1.2. Products with morethan 90% of rare earth oxides can be obtained using this circuit andproduce a concentrated scandium oxide while recovering 80% or more ofthe REEs.

The process of the invention comprises of the two main steps ofsequential precipitation and selective precipitation. An intermediarystep of re-precipitation can further increase RRE concentrations, aswell as improve contaminant metal removal.

For the first step of sequential precipitation, a leachate or aprecipitate obtained from a leachate, is introduced into a solution withan acidic pH value of 2.70 (±0.50 pH units) is placed into a plasticcontainer. The acid used for sequential precipitation is of a differentclass from that used in selective precipitation. The liquid may bestirred or agitated therein, such as through the use of a magneticstirrer to ensure complete dispersion. A base solution, such as sodiumhydroxide or other suitable base of compatible type and concentration,may then be added incrementally, e.g. dropwise, to raise the pH. It isof benefit to continue stirring or agitation to allow for thoroughdispersion so that precipitates formed due to base solutionconcentration gradients are re-dissolved back into the solution.

With the step-wise increase in the solution pH value, precipitates ofiron (III) first appear at around pH 3.20 (±0.20 pH units depending onthe ferric iron concentration) and the majority of the iron is removedat around pH 3.80 (±0.20 pH units). The iron precipitates are thenfiltered out of solution.

Following iron filtration, additional base is added using the sameprocedure, to provide for a gradual increase in the solution pH value toaround 4.80 (±0.20 pH units). At this point, about 85% of the aluminummay now be removed from the liquid as a precipitate. The aluminumprecipitates may then be also removed from the solution. The remainingfiltrate may then serve as a feed solution for the subsequent REEpre-concentration step.

To obtain a REE pre-concentrate, base is added to the remaining filtrateusing the same procedure to elevate the solution pH to about 8.50 (±0.50pH units). By increasing solution pH from 4.80 to 8.50, more than 90% ofREEs are precipitated from the solution. The total REEs content of theprecipitate is typically more than 1% on a dry, whole mass basis. Thefiltrate may then be stored for subsequent REE concentration steps.

For further upgrading purposes, the rare earth pre-concentrate is eitherprocessed using the following procedures of re-precipitation or directlytreated by the selective precipitation process.

Re-precipitation allows an additional, optional step to further increaseREE concentration, for example, when aluminum concentration in thenatural leachate is high (i.e., >500 ppm). The pre-concentrate may beprocessed by mixing the rare earth pre-concentrate with the filtrateobtained after the REE precipitation. Re-dissolution of the precipitateis achieved by adding an acid solution (5M) (i.e., a different class ofacid from that used in sequential precipitation). The acid solution issequentially added in increments, followed by a certain time period(e.g., 5 minutes) of stirring or agitation until the slurry pH isstabilized. This allows for more than 95% of the REEs to be dissolvedback into solution when the pH value is decreased to 2.0 (±0.50 pHunits). A small amount of manganese-rich precipitate cannot be dissolvedand may be then filtered out.

To remove further aluminum from the filtrate, the solution pH isincreased to about pH 4.00 (±0.10 pH units) using a basic solution andthe same incremental adding procedure previously described. Calciumchloride or magnesium chloride may optionally be added into the filtratebefore raising the pH value to promote the aluminum removal efficiency.The aluminum precipitates are removed by filtration. The solution pHvalue of the remaining filtrate is increased to about 9.0 (±0.50 pHunits) by gradual addition of the basic solution (5 M, step 1.2). Inthis case, a rare earth pre-concentrate is obtained with a minor amountof contaminants. The pre-concentrate is used as feed material for thefinal REE concentration step. The total REE content in the precipitateproduced in this step is more than 1% by weight.

After these sequential-concentration and/orre-precipitation-concentration processes, rare earth concentrates ofmore than 90% purity are obtained using the following steps of selectiveprecipitation. Pre-concentrates from the sequential precipitation orfrom the re-precipitation are dissolved into a smaller volume of thefiltrate obtained after precipitation of the REE using the sameprocedure as initially used for re-precipitation, except with a final pHvalue of about 1.2. The solution is filtered and the filtrate stored forsubsequent tests.

The filtrate is then preferably transferred to a separate container andmixed with an acid of a different class from that utilized in sequentialprecipitation and/or re-precipitation, such as e.g., an oxalic acidsolution. The acid solution is optimally stirred to provide for completedispersion. The solution pH value of the suspension after stirring isabout 0.90 (±0.10 pH units).

A strong base solution (e.g. 10 M) is then added incrementally (e.g.0.01 ml per 30 seconds) into the solution. After each addition, thesolution is stirred to ensure that the base is thoroughly dispersed andthe precipitates formed due to base solution concentration gradients arere-dissolved back into solution. The base addition is continued until asolution pH value of about 1.20 (±0.05) pH units is reached.

Stirring may then be maintained for a longer period of time beforeproceeding further, such as for 20 minutes. Solution pH values of thesuspension might deviate slightly from 1.20, in which case base (10 M)or acid (5 M) solution is added in an incremental fashion to maintainthe pH value at about 1.20 (±0.05 pH units).

After stirring, the rare earth precipitates are separated from theresidual liquid using processes such as centrifugation. The finalprecipitates may then be dried, after which rare earth oxide productsare generated by roasting. Rare earth oxide content of the precipitatesis more than 90% and overall rare earth element recovery is more than80%.

1. Examples

A. An example of a possible embodiment for small scale recovery astested on natural leachate samples is as follows:

1.1 Five liter sample of a natural leachate with a solution pH value of2.70 (±0.50 pH units) is placed into a plastic container. The liquid isstirred using a magnetic stirrer at 400 rpm to ensure completedispersion.

1.2 0.5 ml (±0.50 ml) of sodium hydroxide solution (5 M) is added instep wise fashion into the leachate to raise the solution pH value.Magnetic stirring is maintained for 1 minute after each 0.5 ml additionto ensure that the base is thoroughly dispersed and precipitates formeddue to base solution concentration gradients are re-dissolved back intothe solution.1.3 With the step-wise increase in the solution pH value, precipitatesof iron (III) first appear at pH 3.20 (±0.20 pH units depending on theferric iron concentration) and complete removal of iron occurs at pH3.80 (±0.20 pH units). The iron precipitates are filtered out ofsolution using 5 micron (or finer) pore-size filter paper.1.4 After filtration, additional sodium hydroxide solution (5 M) isadded using the same procedure described in Step 1.2 resulting in agradual increase in the solution pH value to 4.80 (±0.20 pH units). Atthis point, about 85% of the aluminum is removed from the liquid as aprecipitate.1.5 The aluminum precipitates are removed from the solution by filteringusing 5 micron (or smaller) pore-size filter paper. The remainingfiltrate serves as a feed solution for the subsequent REEpre-concentration step.1.6 To obtain a REE pre-concentrate, sodium hydroxide solution (5M) isadded to the remaining filtrate using the same procedure described inStep 1.2 to elevate the solution pH to 8.50 (±0.50 pH units). Byincreasing solution pH from 4.80 to 8.50, more than 90% of REEs areprecipitated from the solution. The total REEs content of theprecipitate is typically more than 1% on a dry, whole mass basis. Thefiltrate is stored for subsequent REE concentration steps.

For further upgrading purposes, the rare earth pre-concentrate is eitherprocessed using the following procedures or directly treated by theselective precipitation process. When aluminum concentration in thenatural leachate is high (i.e., >500 ppm), the pre-concentrate isprocessed by:

2.1 Rare earth pre-concentrate obtained in Step 1.6 is mixed with 500 ml(depending on the starting volume) of the filtrate obtained in Step 1.6.Re-dissolution of the precipitate is achieved by adding nitric acidsolution (5M). The acid solution is sequentially added in increments of0.5 ml followed by a certain time period (e.g., 5 minutes) of magneticstirring until the slurry pH is stabilized.2.2 More than 95% of the REEs is dissolved back into solution when thepH value is decreased to 2.0 (±0.50 pH units). A small amount ofmanganese-rich precipitate cannot be dissolved and is filtered out using5 micron (or finer) pore-size filter paper.2.3 To remove aluminum from the filtrate obtained in step 2.2, thesolution pH is increased to pH 4.00 (±0.10 pH units) using 5M sodiumhydroxide solution and the same procedure described in step 1.2. Calciumchloride or magnesium chloride is added into the filtrate before raisingthe pH value to promote the aluminum removal efficiency. The aluminumprecipitates are removed using 5 micron (or finer) pore-size filterpaper.2.4 The solution pH value of the remaining filtrate is increased to 9.0(±0.50 pH units) by gradual addition of sodium hydroxide solution (5 M,step 1.2). In this case, a rare earth pre-concentrate is obtained withminor amount of contaminants. The pre-concentrate is used as feedmaterial for the final REE concentration step. The total REE content inthe precipitate produced in this step is 1.5% to 2.0%. The suspensionwas filtered using 5 micron (or finer) pore-size filter paper.

After the pre-concentration (Step 1.1-1.6) and/or re-pre-concentration(Step 2.1-2.4) processes, rare earth concentrates of more than 90%purity are obtained using selective precipitation as the final process,which is described as follows:

3.1 Pre-concentrates obtained in step 1.6 and/or step 2.4 are dissolvedinto the filtrate obtained in step 1.6 using the same procedure as step2.1 except that the final pH value is 1.2 and the filtrate volume is 100ml. The solution is filtered using 5 micron (or finer) pore-size filterpaper. The filtrate is stored for subsequent tests.3.2 30 ml of the filtrate is transferred to a beaker having a 50 mlmaximum volume. Oxalic acid solution is prepared by dissolving 8 gramsof oxalic acid dehydrate powders (>99% purity) in 50 ml of deionizedwater. Oxalic acid solution of 1.5 ml (optimum amount exists which is afunction of rare earth concentration in solution) is added into thebeaker. The solution is magnetically stirred for 1 minute at 400 rpm forcomplete dispersion. The solution pH value of the suspension afterstirring is about 0.90 (±0.10 pH units).3.3 A sodium hydroxide solution (10 M) is added dropwise (0.01 ml) intothe solution. After each addition, the solution is magnetically stirredfor 30 seconds or longer to ensure that the base is thoroughly dispersedand the precipitates formed due to base solution concentration gradientsare re-dissolved back into solution. The base addition is continueduntil a solution pH value of 1.20 (±0.05) pH units is reached.3.4 Magnetic stirring is maintained for 20 minutes. Solution pH valuesof the suspension might deviate slightly from 1.20, in which case sodiumhydroxide (10 M) or nitric acid (5 M) solution is added in a dropwise(0.01 ml) fashion to maintain the pH value at 1.20 (±0.05 pH units).3.5 After stirring, the rare earth precipitates are separated from theresidual liquid using a centrifuge. The final precipitates are dried inan oven, after which rare earth oxide products are generated by roastingat 800° C. (±100° C.) for 1 hour or longer. Rare earth oxide content ofthe precipitates is more than 90% and overall rare earth elementrecovery is more than 80%.B. Rare Earth Elements Recovery Using Staged Precipitation from aLeachate Generated from Coarse Coal Refuse

Rare earth elements (REEs) are a group of 15 lanthanide elements plusscandium and yttrium, which can be divided into heavy REEs (HREEs: Eu,Gd, Tb, Dy, Ho, Er, Tm, Yb, Lu, Y) and light REEs (LREEs: La, Ce, Pr,Nd, and Sm) based on their locations in periodic table and atomicweights (Binnemans et al., 2013; Massari and Ruberti, 2013). The worlddemand for rare earth elements (REEs) in 2011 was approximately 105,000tones±15% and grew at a rate between 5% and 9% due to their importancein manufacturing of advanced military and renewable energy technologiesas well as many commodity items used by the general public (Alonso etal., 2012; Hatch, 2012). The share of REO demand from wind power,electric vehicles, and NiMH batteries in total clean technologies wasexpected to increase from 11.6%, 50.1%, and 3.4% in 2016 to 13.4%,68.5%, and 10.3% in 2030, but the demand for Nd and Dy from these threefields in 2030 will increase to 199.2% and 268.3% of 2016 level (Klosseket al., 2016; Zhou et al., 2017).

Significant studies have been conducted recently to recover REEs fromcoal and coal related materials to overcome the disadvantages associatedwith commercial rare earth resources (Ayora et al., 2016; Franus et al.,2015; Honaker et al., 2018; Honaker et al., 2017; Honaker et al., 2016a;Honaker et al., 2016b; Honaker et al., 2014; Kashiwakura et al., 2013;Lange et al., 2017; Lin, et al., 2017; Ponou et al., 2016; Rozelle etal., 2016; Zhang et al., 2018; Zhang et al., 2017; Zhang et al., 2015;Ziemkiewicz et al., 2016). In addition to coal, preparation waste, bothcoarse and fine, may also be sources of REEs (Glushkov et al., 2016a,2016b; Vershinina et al., 2016). Coal based materials are normally richin the highly-valued HREEs due to geochemical interactions (Dai et al.,2016; Hower et al., 1999; Seredin and Dai, 2012).

The humic acids associated with coal during the early stages ofcoalification preferentially chelate with the HREEs due to the higherstability of their complexes (Eskenazy, 1987; Pourret et al., 2007;Stern et al., 2007; Wang et al., 2008). REEs in coal mainly exist informs of minerals (either authigenic or detrital), ion-adsorbed and/orion-substituted, and organic bounded forms (Dai and Finkelman, 2018;Seredin and Dai, 2012). Pure monazite particles with >30% REEs wereidentified in fine refuse materials collected from a coal preparationplant processing Fire Clay seam coal (Zhang et al., 2017). Based on themineralogy of the REEs, Honaker et al. (2017) proposed an integratedflowsheet, which combined flotation, hydrophobic-hydrophilic separation(HHS), acid leaching, solvent extraction and selective precipitation, toobtain a final concentrate containing 2% total REEs at a recovery ofaround 50% from coal middlings and refuse materials.

A survey conducted in 2013 of 20 operating coal preparation plantslocated in the northern and central Appalachia coalfields found that thetotal amount of rare earth elements (REEs) contained within the combinedfeed to all 20 plants was nearly 9900 tons annually (Luttrell et al.,2016). Of the total amount of REEs in the plant feed, 63% or 6285 tonsannually existed in the coarse refuse (Honaker et al., 2016a). Honakeret al. (2017) reported that >50% of the REEs associated in wastematerials of three different coal seams (i.e., Fire Clay, West KentuckyNo. 13 and Lower Kittanning) were recovered using sulfuric acid at pH 0,indicating the feasibility of using tank leaching for recovering REEs.To reduce the cost of the leaching process, an alternative method forrecovering REEs from coal refuse is heap leaching utilizing the naturalacid generation capabilities of coal refuse that contains medium-to-highamounts of pyrite (Honaker et al., 2018). The potential benefits includelong term chemical interactions with low capital and operating costs(Pradhan et al., 2008).

Pyrite associated in coal seams can be oxidized into ferric ions in thepresence of oxygen and water along with the generation of a proton (H),which explains the appearance of acid mine drainage (AMD). The REEs aredissolved from the solid material due to the exposure to the naturalleachate, which presents the opportunity for rare earth recovery usinghydrometallurgical techniques such as selective precipitation andsolvent extraction (Da Silva et al., 2009; Migaszewski et al., 2016;Sahoo et al., 2012; Stewart et al., 2017; Sun et al., 2012; Worrall andPearson, 2001a; Worrall and Pearson, 2001b; Zhao et al., 2007;Ziemkiewicz et al., 2016). As shown in FIG. 3, REE concentrationsincreased exponentially from 0.01 ppb to about 2 ppm with a decrease inthe solution pH values from around 8 to 2. A high REE concentrate (e.g.,714 ppm as reported in Ziemkiewicz et al., 2016) was produced in an AMDsludge when using neutralizing agents such as lime to eliminate thenegative impacts of AMD on the environment. Studies by Ayora et al.(2016) and Verplanck et al. (2004) concluded that, instead ofco-precipitating with ferric ions, REEs in AMD tend to be precipitatedat higher pH values (pH>5.1). As such, REE concentration from AMD can berealized by selective precipitation as an alternative to solventextraction to directly recover the REEs from the very dilute AMDsolutions.

A number of studies have been performed previously on metal recoveryfrom acid mine drainages (AMDs) using methods such as precipitation,adsorption, and ion-exchange (Mohan and Chander, 2006; Feng et al.,2000; Wei et al., 2005). Wei et al. (2005) obtained iron and aluminumprecipitates with purity >93.4% and 92.1% from an acid mine drainageusing selective precipitation. Balintova and Petrilakova (2011)recovered 97.16% Fe, 92.9% Al, 95.23% Cu, 88.72% Zn, and 89.49% Mn froman acidic mine drainage in pH ranges <4.05, 4 5.5, 4.49 6.11, 5.5 7.23,and 5.5 9.98, respectively. The redox potential of the AMD was regulatedto control the precipitation behaviors of iron in the solution(Balintova and Petrilakova, 2011; Jenke and Diebold, 1983). Variousneutralizing reagents such as sodium hydroxide, sodium carbonate,ammonia, and lime have been tested. Similar results were reported by Weiet al. (2005) for metal recovery. In addition to the above chemicals,other precipitants such as sodium sulfide and phosphoric acid were alsoutilized to precipitate certain metal ions from AMD. Detailed studies ofthe technical and economic aspects of rare earth recovery from coal minedrainages is limited (Ayora et al., 2016; Ziemkiewicz et al., 2016).

As previously mentioned, the results in the current study are being usedto assist in the design of a heap leach process for recovering rareearth elements. The recovery of REEs and the other valuable elementssuch as cobalt from AMD using selective concentration method is animportant focus of this study, which has received limited attention.

Furthermore, instead of using a single one-step neutralization processfor the acid leachate treatment on mine sites, findings of this studyprovided another approach which also concentrates the associatedvaluable elements. In the study, a natural leachate sample generatedfrom a coal coarse refuse pile of a preparation plant located in westernKentucky was collected together with the solid sample. The waste watersample was generated through long term interaction of clean water withthe coarse refuse, which is similar to the heap leaching process.

Elemental compositions of the collected samples were analyzed toidentify the partitioning behavior of REEs in the liquid relative to thecoarse refuse. To test the feasibility of REE recovery from the naturalleachate, precipitation tests were conducted and precipitates collectedin different pH ranges. Solution chemistry characteristics of REEs inthe leachate were studied by elemental analyses and equilibriumcalculations.

Efficient recovery of REEs from the leachate was finally obtained.

2. Materials and Methods

Coarse refuse generated from the processing of West Kentucky No. 13 seamcoal was identified as a promising feedstock for recovering REEs in aproject funded by the U.S. Department of Energy (Honaker et al., 2016a).Analysis of a representative sample collected from a sweep-belt samplerrevealed ash and total sulfur contents of 83.53% and 6.41%,respectively. The origin of the sulfur was primarily from pyrite.Throughput capacity of the coal preparation plant was 1800 tph whichproduced coarse refuse at a rate of 600 tph. The coarsewaste wastransported and stored in a large refuse pile built on top of afour-foot clay liner located near the processing plant. Natural leachate(waste water) was generated from pyrite oxidization under weatheringeffects at a rate up to approximately 7600 m/day. As required byenvironmental regulations, the pH value of the natural liquid waselevated using lime prior to discharge. A representative liquid samplewas collected from the natural leachate stream at a point locatedupstream of the lime addition for the purposes of the test program.

Upon delivery of the samples to the laboratory, the coarse refuse solidsample was dried under natural atmospheric conditions before beingcrushed and ground to finer than 0.15 mm. A representative sample wascollected from the pulverized material using cone quartering and ariffle splitter. To analyze its elemental composition, the solid samplewas digested using aqua regia and hydrofluoric acid based on the ASTMstandard (ASTM D6357 11). The natural leachate sample was filtered using0.45 m pore size filter papers to remove any suspended particles and thesolution pH value measured to be 2.70. Solution pH was measured using aThermo Scientific Orion Star Pro pH meter with an accuracy of 0.002 pHunit. Metal ion concentration in the solution was analyzed using aninductively coupled plasma optical emission spectrometry (ICP-OES) unit.Ion chromatography (IC) was utilized to measure concentrations of majoranions in the natural leachate.

Staged precipitation tests were performed on the natural leachate sampleof 2 L using an apparatus consisting of a pH meter, magnetic stirrer andvacuum filter. For each stage of precipitation, the solution pH wasgradually increased by adding 2M sodium hydroxide solution.

The pH value was elevated in steps by adding 0.5 mL of the sodiumhydroxide solution prior to each stage. The solution pH before the baseaddition was recorded as the initial pH for this stage of precipitation.To eliminate the potential of localized precipitation due toconcentrated basic solution, the leachate sample was stirred for aperiod of 2 min during and after the base solution addition. Thesolution pH after 2 min of conditioning was recorded as the final pH forthe given precipitation stage. A bulk precipitate was generated due tothe pH increase which was recovered by filtration using 0.45 m pore sizefilter paper. The filtrate was subjected to further pH incrementalincreases followed by filtration to obtain a series of precipitates indifferent pH ranges. The precipitates were dried in an oven for 12 h.

Elemental composition of the precipitate was measured using ICPOES afterre-dissolution of the precipitate following the ASTM D6357 11 method.For each batch of 50 samples, a standard solid sample supplied by theNational Institute of Standards and Technology (NIST) of the U.S.Department of Commerce was digested together with the precipitates. Aliquid standard with known REE contents was also analyzed with eachbatch for quality control of the ICP-OES. Analysis results were acceptedif differences between the measured contents and the values provided bythe supplier were ±10%. Detection limits of the method developed usingthe ICP-OES for REE concentration measurements were: Sc (1.77 ppb); Y(3.72 ppb); La (3.28 ppb); Ce (7.65 ppb); Pr (19.13 ppb); Nd (4.51 ppb);Sm (18.25 ppb); Eu (6.87 ppb); Gd (8.24 ppb); Ho (6.32 ppb); Tb (15.04ppb); Dy (2.30 ppb); Ho (6.32 ppb); Er (2.81 ppb); Tb (15.04 ppb); andLu (7.49 ppb).

Additionally, a fundamental study of the precipitation process for asolution of REEs and other metal ions was conducted based on theelemental compositions of the precipitates and the natural leachate.

Solution equilibrium calculations were performed using the Visual MINTEQ3.1 software which was developed by Jon Petter Gustafsson of KTH,Sweden. The software calculated the equilibrium concentrations andactivities of all the species in the system. Saturation index (SI)values representing the possible solid phases formed in solution atgiven pH values were quantified, which was utilized to determine whetherthe formation of the solid phase was thermodynamically favorable. Thesolution chemistry study provided a fundamental understanding of the REErecovery process from natural coal refuse leachates using the stagedprecipitation method.

Finally, to support the findings of the above studies, sequentialprecipitation tests were performed on three model liquid systems usingconcentrations of specific elements nearly equal to the naturalleachate: liquid 1 (1120 ppm Fe, 1920 ppm SO, 8 ppm La); liquid 2 (540ppm Al, 1920 ppm SO, 8 ppm La); liquid 3 (560 ppm Fe, 270 ppm Al, 1920ppm SO, 8 ppm La). The Fe and Al ion concentrations were varied to keepthe metal to SO ion concentration ratio constant. All the liquids wereprepared by dissolving the required amounts of FeCl.6H O, AlCl.6H O, NaSO, and LaCl.7H O in deionized water. The chemicals were trace metalgrade and purchased from Fisher Scientific. Artificial liquid of 2 Lwere utilized for the tests. The solution pH was continuously increasedby adding sodium hydroxide (2M) and a series of 5 mL supernatants wereobtained from the solution at different pH using a syringe filter toremove the precipitates. The percentage of the elements that wereremoved from the liquid was calculated based on the ICP-OES measurementsof the supernatants and the initial liquid composition.

3. Results and Discussion

3.1. Sample Characterization

Total REE content of the coarse refuse on a dry whole mass basis was 320ppm (Table 1) which was significantly higher than that of the world coalaverage (72.37 ppm) and North American Shale Composite (165.4 ppm)(Gromet et al., 1984; Ketris and Yudovich Ya, 2009). The naturalleachate contained 6.14 ppm of total REEs which was also much higherthan the AMDs at a similar pH value of 2.70 shown in FIG. 3. Ce and Lawere the major REEs in the coarse refuse sample with an overallpercentage of 53% relative to the total REE content. In the naturalleachate sample, yttrium (Y), which is one of the critical REEs (i.e.,Nd, Eu, Tb, Dy, and Y; the others belong to uncritical; defined by Chu,2011), was the dominant REE with a proportion of 28% (FIG. 4) of thetotal REEs. Nearly all of the critical and heavy REEs were concentratedin the natural leachate relative to the coarse refuse. The HREEs/LREEs(H/L) and critical REEs/uncritical REEs (C/UC) ratios of the naturalleachate, which were calculated using mass-based concentrations (ppm),were much higher than that of the coarse refuse (H/L: 1.08 vs 0.21;C/UC: 1.05 vs 0.34). As such, natural acid leaching of the coarse refusevia pyrite oxidization selectively transferred the more valuable andcritical REEs from the solid into the solution. These results indirectlyprovided a reason to pursue the economic recovery of the REEs from coalcoarse refuse using heap leaching processes which would optimize theproduction of the natural leachate.

Previous studies have shown that REEs located in the middle oflanthanide series, i.e., Sm, Eu, Gd, Tb, and Dy, are more likelyenriched in the AMD compared to the REE content of the North AmericanShale Composite (NASC) due to various reasons including: (1) theabundance and distribution of mineral phases containing REEs, (2) thestability of these REE-bearing mineral phases with respect to theaqueous fluids, (3) the chemistry of the aqueous fluids, and (4) theimmobilizing capacity of secondary minerals to REEs (Da Silva et al.,2009; Dai et al., 2013; Stewart et al., 2017; Worrall and Pearson,2001a; Zhao et al., 2007). The REE patterns in the coarse refuse andnatural leachate normalized to the NASC are shown in FIG. 5. Sm, Eu, Gd,Tb, and Dy were more concentrated in the natural leachate compared tothe other REEs when normalized to the NASC, which is in agreement withthe findings reported from previous studies. As such, the naturalleachate generated from the coarse refuse pile shares some similarcharacteristics with AMDs, while the total REE concentration in theleachate was much higher than the AMDs reported in literature (FIG. 3).

3.2. Staged Precipitation

Precipitation behaviors of REEs as a function of the solution pH valueswere investigated through staged precipitation tests. REE contents ofthe precipitates obtained in different pH ranges are shown in Table 2.Thorium contents were also measured for evaluating environmentalimpacts. A precipitate (P8) with 10,763 ppm (1.08%) of total REEs wasobtained in the pH range of 4.85 6.11. H/L and C/UC ratios of theprecipitate were 1.54 and 1.51 (Table 3), respectively, indicatingselective enrichment of the highly-valued REEs in the sample. In morebasic conditions (i.e., pH 6.11 8.55), enrichment of REEs also occurredwith an overall concentration of 2968 ppm. FIG. 6 shows the distributionof total REEs, scandium, thorium, heavy REEs and critical REEs indifferent precipitates calculated based on the rare earth concentrationof the natural leachate (Table 2). Both critical and heavy REEs weremostly recovered in the precipitates produced in the pH ranges 4.85 6.11(74.3% and 74.7%) and 6.11 8.55 (15.4% and 13.5%) with an overallrecovery of >85%, indicating the efficiency of REE recovery from thenatural leachate through staged precipitation.

Enrichment of REEs from coal and coal byproducts have been reportedpreviously. Products containing 1.8%, 0.5%, 0.7%, and 0.13% of totalREEs have been reported by Honaker et al. (2017), Zhang et al. (2017),Zhang et al. (2018), and Lin et al. (2017), respectively, fromcoal-related sources using physical beneficiation methods. However, thegeneration of a REE pre-concentrate from a coal-based source usingstaged precipitation has not been previously reported.

Specific attention needs to be paid to scandium precipitation behaviordue to its significantly higher market value relative to the other REEs.The maximum concentration of scandium occurred in the pH range 4.85 6.11(Table 2), while the majority (37%) was recovered in the pH range 4.674.85 which is reflective of the higher weight percentage (FIG. 6 andTable 3). The proportions of scandium distributed in the lower pHranges, i.e., 3.23 4.08, 4.08 4.55, 4.55 4.64 and 4.64 4.67 were higherthan the other REEs.

Thorium content in the original leachate was low (0.5 ppm) and maximumrecovery in the precipitates occurred in the lower pH range of 3.234.08. Only 10% of the thorium precipitated simultaneously with REEs inthe pH range of 4.85 6.11. As such, thorium was less concentrated in thepre-concentrate compared with REEs. However, thorium existing in thepre-concentrate may be simultaneously enriched with the REEs in thedownstream enrichment process, which may require attention in futurestudies.

Full elemental analyses of the precipitates were performed to evaluatethe selectivity achieved as a function of contaminate cation content andthe potential to recover valuable metals. As shown in Table 4, iron(Fe), aluminum (Al) and magnesium (Mg) were the major components inprecipitates P1-P4, P5-P8 and P9-P12, respectively. The precipitate(P8), which has the highest REE content, contained 1.7% Zn, 1.4% Cu,0.5% Ni and 0.2% Co while precipitate (P9) had 1.5% Ni, 1.3% Zn, 0.5% Coand 0.3% Cu. The above concentrations were close to and/or higher thanthe grades of typical economical ore deposits.

Therefore, a multi-element recovery scheme may be achievable to enhancethe economic value associated with the natural leachate.

Titration of the natural leachate with a strong base such as sodiumhydroxide contributed to the understanding of the buffering reactions inthe aqueous system (Totsche et al., 2003; España et al., 2006).

During the staged precipitation tests, the amount of sodium hydroxideadded and the corresponding solution pH were recorded, which was used todraw a curve similar to a titration curve. Based on the curve (FIG. 7)and the precipitate compositions (Table 4), it was found that twodistinct buffers (i.e., Fe and Al buffering) played significantfunctions in solutions having a pH<4.0 and between 4.0 and 6.0,respectively, corresponding to the two plateaus on the titration curve.The colors of the precipitates shown in FIG. 7 agree with the elementalcompositions. The light brown color of precipitate P8, which containedthe highest amount of REEs, falls in between P6 and P8, which is due tothe enrichment of Al in precipitate P6 and Ni, Mn, and Cu in P9.Aluminum oxides normally appear white while oxides of the other threemetals have a dark color.

3.3. Solution Chemistry Study

Solution equilibrium calculations were conducted in the study using theVisual MINTEQ software to predict precipitation behaviors of REEs andother valuable elements between the leachate and precipitates. Inaddition to equilibrium concentrations and activities of various speciesin solution, the Visual MINTEQ also calculated saturation indexes (SI)to determine whether precipitates are thermodynamically favored toprecipitate or dissolve in solution. The saturation index can berepresented by the following expression:SI=Log IAP−Log K _(ap)where IAP and K represent the ion activity product and solubilityproduct, respectively. IAP equals [A] [B] for a precipitate with A Bformula. SI>0 means the solution is oversaturated and precipitation islikely to occur, while precipitates do not form when SI<0.Concentrations of detectable cations and ions (Table 5) in the leachatemeasured using ICP-OES and IC were utilized as input information forsolution chemistry modelling at pH 2.70. As reported in the previoussection, precipitates were formed as the pH value was increased whichresulted in selective removal of various species from solution. As such,to accurately model the precipitation behavior of the REEs as a functionof solution pH, the input information regarding the concentrations insolution of all species was modified based on the precipitatecompositions listed in Tables 2 and 4.

The saturation index (SI) values of rare earth hydroxides, copper, zinc,iron, and aluminum minerals in equilibrium with the solution underdifferent pH conditions are shown in FIG. 8. The modelling resultsindicate that precipitation of the REEs is thermodynamically favorablein the pH range of 7.0 10.0 which is higher than the range (pH 4.856.11) observed in the staged precipitation tests (FIG. 8(a) and FIG. 6).Fractionation of REEs to solid phases in pH ranges close to the 4.856.11 has been reported previously (Sun et al., 2012; Verplanck et al.,2004; Bau, 1999). Bau (1999) found that REEs are likely to be fractionedto a solid phase in the pH range of 4.6 6.0 due to the formation of ironoxyhydroxide. Verplanck et al. (2004) oxidized six natural AMDs inlaboratory experiments conducted under ambient conditions and found thatREEs initiate partitioning to the solid phase at pH values>5.1 despitethe fact that iron precipitation occurred below pH 5.1. Sun et al.(2012) performed adsorption experiments using acid mine drainage andreported that REEs can be adsorbed onto Mn secondary minerals formed inthe solution in the pH range 4 6. As such, it was concluded that theeffects of secondary minerals formed by the dominant species was thereason for the fractionation of REEs in the lower pH ranges. Based onprevious research, the REE recovery with participates formed underacidic conditions may be caused by a combination of different factors:electrostatic attraction, surface precipitation and ternary complexes.Webster et al. (1998) reported that SO contributes to metal adsorptiononto iron hydroxysulfate surfaces mainly through the formation ofternary complexes between the oxide surface, SO, and the metal ions.

The concentrations of each REE found in the precipitate products listedin the columns of Table 2 were correlated with the concentrations ofeach of the measured non-rare earth elements listed in Table 4 todetermine an overall correlation coefficient. The objective was toquantify the relationship between each REE with the other elementsduring the staged precipitation process. The results listed in Table 6indicated that REEs show a strong positive correlation with Zn, Cu andSi and a medium positive correlation with Al which is reflected by theirrespective high content values in the P8 precipitate product. In the pHrange 4.85-6.11, aluminum hydroxides (e.g., gibbsite) andhydroxysulphates are likely to be the dominant species in theprecipitate as indicated by elemental analyses (Table 4) and saturationindexes (FIG. 8(d)), which agrees with the findings of previous studies(Espana et al., 2006; Totsche et al., 2003). Adsorption of Cu onaluminum hydroxysulfate precipitates has also been reported (Ayora etal., 2016; Rothenhofer et al., 2000). However, in the current study,precipitation of copper is likely in the form of hydroxysulphates, i.e.,antlerite (Cu (OH) SO) and brochantite (Cu (OH) SO), which occur in thepH range of 4.85-6.11 as indicated by the positive SI values (FIG.8(b)).

Silicon in solution occurs largely as undissociated H SiO and is apolymeric colloid at concentrations above its solubility limit (Jenne,1976). As such, precipitation of aluminum, copper and silica whichoccurred in the pH range of 4.85-6.11 explains the enrichment of theother metal ions such as Zn, Mn, and Co. The precipitates formed in anacid mine leachate were poorly crystalized, had very small particlesizes and extremely high specific surface area, which contributed to theadsorption of metal species (Espana et al., 2006; Webster et al., 1998).The adsorption of Zn is indirectly proven by the fact that precipitationof Zn is likely to occur at much higher pH values based on the solutionchemistry modelling results (FIG. 8(b)). The effects of Mn on thefractionation of trace elements have been reported previously, while inthe current study the correlation between Mn and REEs was negligible(Table 6). The function of Mn cannot be ignored given the fact that1.14% of Mn existed in precipitate P8.

Adsorption on the precipitate surfaces via either surface precipitationor ternary complexation is controlled by the stability of the complexesformed between the metal ions and SO, which can be investigated by thespeciation distribution. Proportions of Sc, Y, Zn, Co and Ni thatoccurred as free species are listed in FIG. 9. Nearly 100% of thecomplexed species of the above elements occurred as sulfates. Thepreferential fractionation of Sc to the solid phase that occurred in thelower pH ranges may be explained by that fact that Sc is more likelycombined with sulfate anions. The highest concentrations of Co and Nioccurred in the higher pH range 6.11-8.55 instead of 4.85-6.11, whichcan also be attributed to the fact that (i) the two elements have loweraffinity for SO and (ii) the fraction of sulfate complexes increaseswith an evaluation in the solution pH values. Ochreous oxideprecipitates (i.e., poorly crystallized oxy hydroxyl sulfates of Fe)which typically comprise weakly crystallized goethite, jarosite, ferriteand schwertmannite were major components of the precipitates obtained atpH values<4.55 as indicated by the elemental composition (Table 4) andsaturation index (FIG. 8(c)) (Jonsson et al., 2006; Verplanck et al.,2004). The current study indicates that the effects of iron precipitateson the fractionation of REEs were minor, which may be due to lowerstability of the rare earth sulfate complexes at lower pH values and/orcompetitive adsorption between the REEs and the other species such asAl, As and Pb.

3.4. Model System Study

To investigate the interactions among Fe, Al, and La in solutionscontaining SO, sequential precipitation tests were performed on thethree model liquid systems, i.e., Liquid 1 (1120 ppm Fe, 1920 ppm SO, 8ppm La), Liquid 2 (540 ppm Al, 1920 ppm SO, 8 ppm La) and Liquid 3 (560ppm Fe, 270 ppm Al, 1920 ppm SO, 8 ppm La). The Fe and Al ionconcentrations were varied to keep the metal to SO ion concentrationratio constant.

Percentage removal of the elements from the three liquids are shown inFIG. 10.>95% of the iron and aluminum were removed at pH values 3.0 and4.5, respectively, for all three liquids. Removal of iron in liquid 3was lower than liquid 1 for solution pH values<3.0, which may beexplained by the lower iron concentration in liquid 3 (560 ppm vs 1120ppm). However, aluminum removal in liquid 3 was higher than liquid 2 forpH values less than about 3.8 despite the fact that its concentration inliquid 3 was only half of liquid 2. In another word, precipitation ofiron contributed to the fractionation of aluminum to the solid phase.

Unlike iron and aluminum, which were removed in narrow pH ranges of 2.0to 3.0 and 3.5 to 4.5, respectively, the removal of lanthanum increasedgradually as the solution pH value was elevated ((b)). As FIG. 10 such,instead of co-precipitation, lanthanum was removed from the liquids morelikely through adsorption onto iron and aluminum hydroxides and/orhydroxysulfates surfaces, which agrees with the solution chemistry studyof the real leachate. As shown in FIG. 10(b), the removal of lanthanumfrom the solution containing only Fe and SO (Liquid 1) at pH3.5 wasaround 40%. However, in Liquid 2 which contained Al and no Fe, there wasno removal of lanthanum or aluminum at pH 3.5 indicating that noparticipation occurred in the system. When both Fe and Al were presentin solution with SO and La, nearly 10% of the La ad Al was removed whichwas most likely due to competitive adsorption onto the ironhydroxysulphate precipitates. As the pH was increased above 3.5, bothiron and aluminum hydroxysulphates precipitated which resulted innon-competitive adsorption of La onto the precipitate surfaces and 100%removal at a pH of around 6.5. The findings of the model system studyprovided an understanding for the selectivity achieved when treating thenatural leachate, i.e., 1. The presence of aluminum sulfate limited theamount of rare earth elements that adsorbed onto the ironhydroxysulphate precipitate surfaces at pH values<4.85 and; 2. Both ironand aluminum hydroxysulphates precipitated at pH values<4.85 therebyeliminating nearly 100% of the iron and the majority of the aluminum.

3.5. Upgrading of the REE-Enriched Precipitate

The precipitate fractions enriched in REEs requires further upgrading tobe commercially marketable as a mixed concentrate. Precipitate materialidentified as P8 and P9 in Table 3 were mixed together and re-dissolvedusing 10M HNO. The indissoluble material was removed using a 0.45 m poresize filter paper.>95% of REEs associated in the pre-concentrate (i.e.,P8 and P9) were recovered into the solution. As shown in Table 7, thetotal REE content in solution was 52 ppm of which Y and Nd accounted for50% of the total. Aluminum and magnesium were the primary contaminants.

After the dissolution step, oxalic acid was added to the pre-concentratesolution to selectively precipitation the REEs which resulted in theproduction of a high-grade mixed REE product. The oxalate precipitateswere dried in an oven at 60° C. for 12 h and then roasted at 750° C. for2 h. Under a given set of conditions (i.e., pH 1.20; aging time 30 min;oxalic acid dosage 0.02M), a final product containing 94% rare earthoxides (REO) was obtained with the REE distribution shown in Table 8.Approximately 30% of the mixed REO product was Nd O, Pr O, Dy O whichare compounds commonly used in the manufacturing of permanent magnets.The overall REE recovery from the pre-concentrated solution was 78%. Thefinal product contained about 0.5% Th and 0.2% U which is sufficientlyhigh to require another process step. Several alternatives exist toreduce the concentration of the radioactive elements including causticconversion, solvent extraction and ion adsorption technologies (Zhu etal., 2015).

3.6. Economic Assessment

A typical application of the staged precipitation process for REErecovery is the treatment of acidic water generated from the oxidationof pyrite at coal mining operations. For the source of the acidic waterin this study, the volumetric flow rate of acidic water is approximately500 m/h which contains 6.14 ppm of total REEs on a weight basis.

Assuming that a REE recovery plant would operate 24 h/day and 365 daysannually due to the continuous flow, the annual production of mixed REEconcentrate would be approximately 27 t using a recovery factor of 80%based on experimental results. For a typical 10-year life for theequipment utilized in the recovery process, the total amount of mixedREE concentrate generated by a selective precipitation plant would bearound 270 t.

The economic value of the mixed REE concentrate produced from treating anatural leachate is difficult to assess due to the volatility of themarket values and the associated costs of the downstream process neededto produce individual rare earth concentrates. However, Seredin and Dai(2012) introduced an expression that quantifies an outlook coefficient(C) based on the total percent content of critical rare earth elementsin a given material relative to the total REE content (REY), i.e.:

${REY}_{{def},{rel}} = {{100\% \times \frac{{Nd} + {Eu} + {Tb} + {Dy} + Y}{{\sum\;{REY}};}C_{outl}} = {\left( {\left( {{Nd} + {Eu} + {Tb} + {Dy} + Y} \right)/{\sum\;{REY}}} \right)/\left( {\left( {{Ce} + {Ho} + {Tm} + {Yb} + {Lu}} \right)/{\sum\;{REY}}} \right)}}$

The outlook coefficient value of 3.85 for the natural leachate sourceused in this study compares favorably with well-known REE minesextracting monazite and bastnaesite as shown in Table 9. The value ofthe natural leachate is realized by the amount of neodymium anddysprosium which represent nearly 20% of the total REE content. Inaddition, nearly 9% of the total REE content is scandium which has thehighest market value of all the REEs. The xenotime sources listed inTable 9 have higher outlook coefficient values due to elevated yttriumcontents.

The cost of the selective precipitation process when treating thenatural leachate is significantly enhanced by the fact that treatment ofthe acidic water source is mandated by regulatory agencies. A review byJohnson and Hallberg (2005) states that chemical treatment using analkaline solution is the most commonly used technique for watertreatment at a mine site. As such, the chemical cost of elevating thesolution pH value through the stage precipitation process does not needto be included in the overall cost. The chemical costs that are requiredinclude nitric acid to re-dissolve the REE pre-concentrate(precipitates) back into solution (10 t/m of leachate) and oxalic acidneeded to selectively precipitate the REEs from the final leachate (10t/m of leachate). Additional chemicals will be needed to addressenvironmental issues such as thorium and uranium.

A preliminary cost assessment was conducted based on the treatment of500 m/h of natural leachate using the project cost estimation guidelinesdescribed by Mular (2002). The project cost sheet presented in Table 10includes the capital cost for equipment that essentially consists of aseries of tanks and thickeners needed to achieve the necessary feedconditioning, pH adjustment and precipitate concentration.

Total equipment cost is multiplied by a series of factors to estimatethe costs of ancillary items. Indirect costs are expressed as apercentage of the total indirect costs. The manpower cost is based onone supervisor and one individual operating the system each shift. Theplant will operate 3 shifts daily and 365 days annually. The totalcapital and operating cost needed to recover 270 t of mixed REEconcentrate over a 10-year period was estimated to be around $9.2million. As such, the effective cost of producing 1 kg of mixedconcentrate from the given source using the process described in thispublication is approximately $33.9.

4. Conclusions

Staged precipitation tests were conducted in the current study for anatural leachate collected from a coal coarse refuse pile. The naturalleachate contained 6.14 ppm of total REEs due to dissolution of thesolid waste material by the acid generated from pyrite oxidization. Aprecipitate containing 1.1% total REEs, of which 64% are consideredcritical REEs, was obtained from the leachate in the pH range of4.85-6.11. The precipitate also contained 18.4% Al, 1.7% Zn, 1.4% Cu,1.14% Mn, 0.5% Ni and 0.2% Co, indicating the potential of additionaladded value from metals other than REEs. A mixed product containing 94%rare earth oxides was obtained by dissolution of the precipitatesenriched in REEs followed by selective precipitation using oxalic acid.The final product content was especially high in yttrium, neodymium,samarium, gadolinium, and dysprosium oxide indicating the potential forsignificant economic value.

Precipitation characteristics of the REEs and other metal ions wereinvestigated through solution chemistry calculations and modelling. Thesolution chemistry study indicated that enrichment of REEs and the othervaluable elements in the precipitates was due to the adsorption effectsof hydroxides and hydroxysulfates of Al, Si and Cu. The sequentialprecipitation tests performed on model systems indicated thatcompetitive adsorption on the iron precipitate surfaces existed betweenAl and the REEs, which explained the selectivity realized in the stagedprecipitation process when treating the natural leachate.

TABLE 1 Rare earth contents (ppm, on a dry whole mass basis) in thecoarse refuse and leachate samples. Sample Sc Y La Ce Pr Nd Sm Eu Coarse19.10 28.50 51.20 119.40 19.20 45.00 10.80 1.80 refuse Leachate 0.541.75 0.32 0.60 0.22 0.89 0.41 0.08 Sample Gd Tb Dy Ho Er Tm Yb Lu Coarse9.00 0.80 4.50 1.20 4.70 0.55 3.40 0.48 refuse Leachate 0.53 0.09 0.350.05 0.14 0.03 0.12 0.02

TABLE 2 Individual and total rare earth concentrations on a dry, wholemass basis (ppm) for precipitates obtained from the staged precipitationtests. Sample pH_(initial) pH_(final) Sc Y La Ce Pr Nd Sm Eu P1 2.7 2.941.3 4.7 1.6 12.2 50.6 0.0 20.1 0.4 P2 2.94 3.23 1.5 4.6 1.5 12.6 51.30.0 20.1 0.4 P3 3.23 4.08 22.0 28.0 4.5 34.8 56.8 24.5 32.2 2.7 P4 4.084.55 66.9 14.6 8.5 30.4 8.1 9.1 5.0 0.8 P5 4.55 4.64 78.0 15.3 4.0 12.82.0 5.5 2.7 0.6 P6 4.64 4.67 80.4 14.4 2.1 10.2 1.4 4.0 1.0 0.5 P7 4.674.85 116.4 30.9 1.8 11.2 1.5 6.9 2.5 0.9 P8 4.85 6.11 235.6 3785.3 210.31057.6 339.0 1693.6 702.1 169.9 P9 6.11 8.55 41.8 783.5 200.0 524.2109.6 566.4 179.8 31.3 P10 8.55 10 14.5 267.7 69.6 172.6 37.6 191.7 58.69.7 P11 10 10.83 1.2 13.4 2.4 7.1 2.4 9.9 3.7 0.5 P12 10.83 12 2.0 2.80.0 0.0 1.7 2.2 1.2 0.2 Sample Gd Tb Dy Ho Er Tm Yb Lu Th P1 2.3 7.114.6 0.0 0.0 0.0 3.9 8.1 6.2 P2 2.4 7.0 15.0 0.0 0.0 0.0 3.9 8.1 11.5 P313.7 8.8 23.5 1.0 2.9 0.5 7.5 7.8 121.1 P4 5.5 1.4 5.4 1.1 1.7 0.6 2.10.4 101.4 P5 4.0 0.7 3.5 0.9 1.6 0.4 1.8 0.0 48.6 P6 3.4 0.5 3.2 0.8 1.50.4 1.8 0.0 37.4 P7 5.8 1.1 7.3 1.3 3.6 0.8 4.5 0.2 18.0 P8 891.3 132.5789.8 129.9 321.5 52.4 221.1 35.3 69.2 P9 216.3 21.5 148.8 26.0 62.410.5 40.0 5.5 39.1 P10 72.6 2.2 49.3 9.0 20.7 3.4 13.1 6.3 15.4 P11 3.60.3 2.6 0.5 1.0 0.1 0.6 0.0 0.5 P12 1.3 0.0 0.6 0.3 0.2 0.0 0.1 0.0 0.0

TABLE 3 Precipitate weight (%), total REEs (ppm, on a dry whole massbasis), H/L and C/UC ratios, and total REE recovery (%) for theprecipitates obtained in different pH ranges. Re- Weight REEs coverySample pH_(initial) pH_(final) (%) (ppm) H/L C/UC (%) P1 2.7 2.94 21.65127.0 0.48 0.27 3.29 P2 2.94 3.23 16.94 128.4 0.48 0.26 2.60 P3 3.234.08 9.64 271.2 0.55 0.47 3.13 P4 4.08 4.55 4.09 161.8 0.26 0.24 0.79 P54.55 4.64 6.06 133.8 0.27 0.24 0.97 P6 4.64 4.67 5.04 125.6 0.27 0.220.76 P7 4.67 4.85 14.55 196.9 0.40 0.31 3.43 P8 4.85 6.11 5.08 10,763.21.54 1.51 65.40 P9 6.11 8.55 4.46 2967.5 0.83 1.07 15.83 P10 8.55 0 2.79998.4 0.83 1.07 3.33 P11 10 10.83 7.31 49.3 0.85 1.15 0.43 P12 10.63 122.42 12.7 0.78 0.82 0.04

TABLE 4 Major and trace elmental contects, (ppm, if not mentioned) ofthe precipitates collected from the staged precipitation test. SamplepH_(initial) pH_(final) Fe (%) Al (%) Na (%) Mg (%) Ca (%) Mn (%) As P12.7 2.94 40.6 0.19 0.10 0.08 0.07 0.05 579 P2 2.94 3.23 41.3 0.24 0.110.07 0.07 0.08 154 P3 3.23 4.08 39.5 1.50 0.19 0.07 0.07 0.15 49 P4 4.084.55 44.0 18.1 0.45 0.11 0.10 0.18 0 P5 4.55 4.64 0.51 19.7 0.51 0.120.11 0.11 0 P6 4.64 4.67 0.27 20.0 0.49 0.10 0.09 0.13 0 P7 4.67 4.850.07 20.2 1.00 0.10 0.09 0.17 0 P8 4.85 6.11 0.02 18.4 1.79 0.74 0.311.14 0 P9 6.11 8.55 0.04 4.28 2.12 15.4 0.34 0 0 P10 8.55 10 0.03 1.532.12 20.6 2.06 0 0 P11 10 10.83 0.02 0.15 1.85 19.7 0.17 6.88 0 P1210.83 12 0.03 0.33 2.11 32.3 0.11 0.25 0 Sample Si Pb Co Zn Cd Cr Cu NiMo Sn B P1 332 150 53 15 13 4 0 0 32 43 14 P2 363 129 57 15 14 31 0 0 1426 17 P3 322 138 58 17 11 247 192 0 16 40 15 P4 242 0 0 34 0 130 107 7616 61 12 P5 138 0 0 56 0 59 87 86 0 25 12 P6 158 0 0 45 0 55 100 90 1315 21 P7 142 0 0 65 0 16 309 149 11 40 13 P8 2040 15 1875 16,904 0 2714,012 5487 12 54 21 P9 168 0 5121 13,169 170 96 3226 15,170 0 55 11 P10196 0 1649 3912 38 35 1043 4419 9 19 29 P11 343 0 36 176 0 0 22 152 6 1636 P12 313 0 0 0 0 0 0 0 0 58 20

TABLE 5 Elemental concentrations (ppm) of major species in solution. FeAl Mg ca Mn Na Zn Si Ni Cu 1524 697 457 293 85.3 24.6 13.1 10.3 8.6 8.7Co Cr As Sb Pb Sn U SO₄ ²⁻ Cl⁻ 3.2 0.2 0.2 0.3 0.3 0.2 0.2 11,420 14.4

TABLE 6 Pearson correlation coefficients of REEs with the other elementscalculated based on the elemental compositions of the precipitates(P1-P12). Coefficient Fe Al Mg Ca Mn As Si Pb Co Zn Cd Cr Cu Ni Mo Sn BSc 0.34 0.77 0.43 0.11 0.12 0.32 0.74 0.36 0.17 0.61 0.15 0.05 0.81 0.180.05 0.26 0.17 Y 0.27 0.28 0.12 0.08 0.03 0.16 0.95 0.16 0.43 0.88 0.070.11 1.00 0.43 0.04 0.34 0.07 La 0.56 0.09 0.10 0.26 0.07 0.22 0.62 0.260.87 0.99 0.63 0.01 0.82 0.87 0.25 0.38 0.04 Ce 0.31 0.21 0.04 0.16 0.020.19 0.85 0.20 0.66 0.97 0.34 0.04 0.97 0.66 0.13 0.39 0.01 Pr 0.11 0.130.16 0.06 0.04 0.02 0.93 0.05 0.49 0.88 0.17 0.00 0.97 0.49 0.08 0.370.01 Nd 0.30 0.24 0.07 0.12 0.01 0.18 0.91 0.18 0.54 0.93 0.20 0.08 0.990.54 0.09 0.36 0.05 Sm 0.25 0.24 0.12 0.08 0.01 0.14 0.95 0.12 0.47 0.890.12 0.08 1.00 0.47 0.03 0.36 0.05 Eu 0.26 0.29 0.13 0.06 0.03 0.15 0.960.14 0.41 0.86 0.05 0.10 1.00 0.41 0.03 0.34 0.07 Gd 0.28 0.27 0.11 0.090.02 0.16 0.94 0.16 0.46 0.89 0.11 0.10 1.00 0.46 0.05 0.35 0.06 Tb 0.180.26 0.19 0.00 0.02 0.09 0.97 0.06 0.36 0.83 0.01 0.08 0.99 0.36 0.040.35 0.04 Dy 0.25 0.27 0.14 0.06 0.02 0.14 0.96 0.12 0.41 0.86 0.05 0.101.00 0.41 0.01 0.34 0.07 Ho 0.27 0.29 0.12 0.07 0.03 0.16 0.95 0.16 0.420.87 0.06 0.11 1.00 0.42 0.04 0.34 0.07 Er 0.27 0.29 0.13 0.07 0.03 0.160.96 0.15 0.42 0.87 0.06 0.11 1.00 0.42 0.03 0.34 0.07 Tm 0.27 0.29 0.130.07 0.02 0.16 0.95 0.16 0.42 0.87 0.06 0.10 1.00 0.42 0.03 0.35 0.07 Yb0.25 0.29 0.15 0.06 0.02 0.14 0.96 0.12 0.40 0.86 0.04 0.10 1.00 0.400.01 0.34 0.06 Lu 0.00 0.08 0.23 0.10 0.06 0.09 0.95 0.19 0.32 0.77 0.000.03 0.93 0.31 0.22 0.29 0.05

TABLE 7 Elemental composition (ppm) of the pre-concentrated solution. ScY La Ce Pr Nd Sm Eu Gd Tb Dy 1.1 17.6 1.5 5.7 1.7 8.6 3.4 0.8 4.2 0.63.6 Ho Er Tm Yb Lu Fe Al Mg Ca Mn 0.6 1.5 0.2 1.0 0.2 2.0 791 543 23 36

TABLE 8 Rare earth oxides contents (%) of the product obtained in thestudy. Sc₂O₃ Y₂O₃ La₂O₃ CeO₃ Pr₆O₁₁ Nd₂O₃ Sm₂O₃ Eu₂O₃ 0.13 16.64 6.0420.33 4.77 20.42 7.16 1.60 Gd₂O₃ Tb₂O₃ Dy₂O₃ Ho₂O₃ Er₂O₃ Tm₂O₃ Yb₂O₃Lu₂O₃ 7.61 1.02 4.92 0.76 1.45 0.19 0.92 0.16

TABLE 9 REY_(def, rel) and C_(outl) of the natural leachate and someother resources; non-coal based data obtained from Chen (2011). SourcesREY_(def, rel) C_(outl) Mountain Pass; bastnaesite 11.4 0.23 Bayan Obo;bastnaesite 19.0 0.38 Green Cove Spring; monazite 21.7 0.49 NaturalCoal-Based Leachate 56.4 3.85 Lehat; xenotime 72.9 5.13 Longnam;xenotime 77.2 13.73 Note: Non-coal based data obtained from Chen (2011).

TABLE 10 Economic assessment of a plant for recovering REEs from thenatural leachate (plant life: 10 years; volumetric throughput capacity:500 m³/h; REEs concentration in leachate 7 ppm; 80% of total REErecovery). Unit cost Number Size (S) required Cost (S) Fixed equipmentMix tank 90 m³ 64,942 4 259,767 Mix tank 10 m³ 17,000 2 34,000Agitator + motor 25 hp 22,173 5 110,863 316 stainless Agitator + motor 5hp 7000 5 35,000 316 stainless Makeup tanks 30 m³ 70,457 6 422,741Thickeners 250 m³ 130,942 3 392,825 Thickeners 30 m³ 15,713 2 31,426Plate and frame 5 m² 36,180 1 36,180 filter Pumps centrifugal 4 in. inoutlet 14,181 10 141,814 Roasting ovens 0.5 m³ 5562 4 22,248 Fixedequipment 1,464,616 subtotal Ancillary Piping 10% of fixed equipmentcost 146,461 Electrical 10% of fixed equipment cost 146,461 Controls 10%of fixed equipment cost 146,461 Ventilation + 5% of fixed equipment cost73,231 HVAC Plumbing 5% of fixed equipment cost 73,231 Ancillary 585,847sub total Facilities Building 5000 m² 180 s/m² 900,000 Total 2,950,463Indirect (Basis) Engineering 10% of the total 295,046 Contingency 20% ofthe total 590,092 Chemical cost Consumption Unit Cost (S/ton) Cost (S)HNO₃ 5 × 10⁻⁶ ton/m² 300 657,000 of leachate Oxalic acid 5 × 10⁻⁵ ton/m³400 876,000 of leachate Other 10% of HNO₃ and oxalic acid 153,300 totalcost Total cost minus labor 5,521,901 Labor cost 3,640,000 Total cost9,161,901 Cost to produce 1 kg of rare earth combined oxides 33.9 (>94%purity) Note: Equipment costs obtained from http://www.matche.com andchemical costs from https://www.allbaba.com/.

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The foregoing has been presented for purposes of illustration anddescription. It is not intended to be exhaustive or to limit theembodiments to the precise form disclosed. Obvious modifications andvariations are possible in light of the above teachings. All suchmodifications and variations are within the scope of the appended claimswhen interpreted in accordance with the breadth to which they arefairly, legally and equitably entitled. All documents referenced hereinincluding patents, patent applications and journal articles and herebyincorporated by reference in their entirety.

We claim:
 1. A method for concentrating rare earth materials from a feedstock comprised of coal leachate and/or precipitate obtained therefrom,the method comprising: a) mixing the feed stock with an acid of a firstclass in a solution; b) increasing the pH of the solution incrementallyto precipitate iron and aluminum and removing such precipitates from thesolution; c) further increasing the pH incrementally to precipitate rareearth metals and separating such from the resulting filtrate; d)re-dissolving the precipitated rare earth metals in a smaller volume ofthe filtrate; e) transferring the re-dissolved filtrate to an acidsolution of a second class; and f) incrementally increasing the pH toselectively precipitate rare earth metals.
 2. The method of claim 1,further comprising a re-precipitation step between c) and d), the stepcomprising re-dissolving the rare earth metals in the filtrate bylowering the pH.
 3. A method for concentrating rare earth materials froma feed stock, the method comprising: a) mixing the feed stock with anacid in a solution; b) increasing the pH of the solution incrementallyto precipitate iron and aluminum and removing the precipitated iron andaluminum from the solution; c) further increasing the pH incrementallyto precipitate rare earth metals and separating the precipitated rareearth metals from a resulting filtrate; d) re-dissolving theprecipitated rare earth metals in a smaller volume of the filtrate; e)transferring the re-dissolved filtrate to an acid solution; and f)incrementally increasing the pH to selectively precipitate rare earthmetals.
 4. A method for concentrating rare earth materials from a feedstock, the method comprising: a) mixing the feed stock with an acid in afirst solution; b) increasing the pH of the solution incrementally toprecipitate contaminants and removing the precipitated contaminants fromthe solution; c) further increasing the pH incrementally to precipitaterare earth metals; d) re-dissolving the precipitated rare earth metalsin the filtrate by lowering the pH; e) incrementally increasing the pHto gradually precipitate and remove additional contaminates and finallyprecipitate a rare earth metal concentrated re-precipitate; f)re-dissolving the rare earth metal concentrate re-precipitate in asmaller volume of the filtrate; g) acidifying the smaller volume of thefiltrate including the re-dissolved rare earth metal concentratere-precipitate; and h) incrementally increasing the pH to selectivelyprecipitate the rare earth metals.
 5. The method of claim 4, includingusing a different acid to re-dissolve the rare earth metals in thefiltrate than used when mixing the feed stock.
 6. The method of claim 5,including adding CaCl₂, MgCl₂ or CaCl₂ and MgCl₂ to the smaller volumeof the filtrate prior to step e).
 7. The method of claim 4, includingadding CaCl₂, MgCl₂ or CaCl₂ and MgCl₂ to the smaller volume of thefiltrate prior to step e).
 8. The method of claim 4, further includingusing coal leachate, a coal leachate precipitate or coal leachate and acoal leachate precipitate as the feed stock.
 9. The method of claim 4,further including using a first class of acid to mix with the feedstockin the solution and using a second, different class of acid tore-dissolve the rare earth metal concentrate re-precipitate in thesmaller volume of the filtrate.